Hydrometallurgical treatment process for extraction of platinum group metals obviating the matte smelting process

ABSTRACT

A hydrometallurgical treatment process for extracting platinum group metals from a flotation concentrate in which the invention revolves around obviating the matte smelting and granulating process. Instead the concentrate is submitted to pressure leaching, oxidative or reductive roasting and final recovery by means of ion exchange adsorption. Roasting is applied in order to convert the platinum group metals to a form that dissolves in chlorine/HCL and a chlorine/HCL leach that renders the platinum group metals in solution.

BACKGROUND TO THE INVENTION

This invention relates to the hydrometallurgical treatment process forextracting platinum group metals from a flotation concentrate.

Conventionally, platinum group metals are extracted from a flotationconcentrate in a matte smelting and converting process followed byfurther refining for the extraction of the platinum group metals.

SUMMARY OF THE INVENTION

According to the invention there is provided a hydrometallurgicaltreatment process for extracting platinum group metals from a flotationconcentrate comprising the steps of:

leaching of the flotation concentrate to dissolve base metal sulphidesin the flotation concentrate so as to form a filtrate and a residue;

separation of the filtrate from the residue;

roasting the residue to form a calcine; and

chlorination of the calcine to dissolve the platinum group metals intosolution.

Typically, the process includes the additional steps of:

adsorption of the platinum group metals onto an ion exchange resin; and

recovery of the platinum group metals from the ion exchange resin.

Preferably, the roasting step involves oxidation or reduction, morepreferably oxidation at up to 1000° C.

Typically, the method includes the step of recovering Osmium from theoff-gas from the roasting step.

The chlorination step preferably comprises countercurrent chlorinationof the calcine at approximately 80° C. and 3.5N HCl.

The separation step typically comprises filtration followed by theadditional steps of neutralisation of the filtrate; precipitation ofbase metal sulphides and flotation of precipitated sulphides into aconcentrate.

The step involving adsorption of the platinum group metals onto an ionexchange resin may be followed by:

desorption of the platinum group metals from the resin with thiourea atapproximately 80° C. followed by water washing of the stripped resin;and/or

precipitation of the platinum group metals from the eluate with causticsolution.

Various embodiments of the invention are described in detail in thefollowing passages of the specification which refer to the accompanyingdrawings. The drawings, however, are merely illustrative of how theinvention might be put into effect, so that the specific form andarrangement of the features shown is not to be understood as limiting onthe invention.

BRIEF DESCRIPTION OF THE ACCOMPANYING DRAWINGS

FIG. 1 is a diagrammatic flow sheet of a first embodiment of thehydrometallurgical extraction process of the invention;

FIG. 2 is a table which sets out the composition of a flotationconcentrate which is used to describe the first embodiment of the methodof the invention;

FIG. 3 and FIG. 3A comprises two tables setting out the results achievedin experimental work on the autoclave oxidative leaching of a sample offlotation concentrate; and

FIG. 4 is a diagrammatic flow sheet of a second embodiment of thehydrometallurgical extraction process of the invention.

DESCRIPTION OF AN EMBODIMENT

FIG. 1 of the accompanying drawing depicts diagrammatically a firstembodiment of the hydrometallurgical treatment process according to theinvention for extracting platinum group metals from a flotationconcentrate. In broad outline the proposed process comprises thefollowing unit operations:

autoclave oxidative leaching of the concentrate to dissolve the basemetal sulphides;

filtration of the oxidised slurry;

neutralisation of the filtrate and precipitation of base metal sulphideswith lime/sulphur, followed by flotation of the precipitated sulphidesinto a concentrate;

oxidative roasting of the residue;

scrubbing of the off-gas from the roaster for Os recovery;

countercurrent chlorination of the calcine which is the product of theroasting step;

cooling and filtration of the chlorinated slurry with washing of thefilter cake;

disposal of the washed residue;

adsorption of the platinum group metals from the filtrate onto an ionexchange resin;

desorption of the platinum group metals from the resin with thioureafollowed by water washing of the stripped resin;

precipitation of platinum group metals from the eluate with causticsolution;

thickening and filtration of the platinum group metal precipitate; and

removal of iron from the resin washing solution by solvent extractionwith a tertiary amine, the iron and other base metals being strippedfrom the extractant with water and then precipitated with soda ash.

The process will now be described in greater detail with reference tothe accompanying drawings and tables.

In order to illustrate the first embodiment of the invention a flotationconcentrate is used having a composition as is set out in FIG. 2. Theplatinum group metal flotation concentrate is introduced into theprocess as feed 1. The feed prepared as a slurry 2 is subjected toautoclave leaching 3 in order to dissolve, at least partially, basemetals such as Ni, Cu, Co and Fe. This is done prior to the leaching ofthe platinum group metals from the concentrate so as to remove the basemetals from the process and thereby simplify the recovery of theplatinum group metals.

Any iron which remains in the solid phase, mainly in the hydrated form,would have a negative influence on the results of further stages such ascalcination, chlorination or adsorption. A process which may beimplemented to assist with the removal of iron at the initial stage isto pre-treat the initial concentrate with sulphuric acid in an autoclavewithout the presence of an oxidiser such as oxygen. Without the properlychosen process perameters sulfide iron, present in the form ofpyrrhotite, pentlanddite and chalcopyrite, decompose and transfer to thesolution in the form of FeSO₄.

The dissolution of the base metals is standard technology and istypically done by oxidation under pressure in an autoclave, at an oxygenpressure of 1.0 MPa, a liquid to solid ratio in the flotation slurry of3 and a temperature of 150° C. with a residence time of 1.5 hours.

Autoclave leaching also has the advantage of removing sulphur which ispresent in the concentrate. This is beneficial as it leads to reducedSO₂ handling in the subsequent roasting stage. Through experimental workit was found that the autoclave leaching of a platinum group metalflotation concentrate having a composition as is depicted in FIG. 2 andapplying the aforementioned conditions results in desirable recovery ofsulfides with a transfer of 93 to 96% of nickel and more than 70% ofcopper to the solution. Transition to the solution among platinum metalsis found to be low, in the region of 2 to 2.5% of the quantity of metalin the initial concentrate. It was found that the degree of Pt and Pddissolving was less than 0.5%.

FIG. 3 sets out the results that were achieved in the autoclaveoxidative leaching of a concentrate sample having a chemical compositionset out in FIG. 2. These experiments in leaching were carried out in 1and 3 liter capacity autoclaves at a temperature of 150° C., partialoxygen pressure of 1 MPa, rotation speed of a turbine mixer @ 2800min⁻¹, a liquids to solids ratio of between 2 and 3 and a processduration of 40 to 120 minutes. The results of the experimental work arepresented in the table of FIG. 2. In this table only the consumption ofNi and Cu into solution are recorded.

From the results set out in table 2 of FIG. 2 and a series of otherexperimental work that was conducted on various concentrate samples thefollowing mode of oxidizing leach was found to be desirable foroxidizing leaching of flotation concentrates with relatively highsulphur content:

temperature 150° C.

oxygen partial pressure 1 MPa

process duration 60-80 minutes

liquids to solids ratio 3.

It is important that in the base metal removal stage 3 the quantities ofplatinum group metals that are dissolved are kept to a minimum. Underthe conditions specified above it has been found there is negligibledissolution of platinum group metals.

After base metal dissolution the resultant slurry is filtered 4, withthe filtrate being processed to recover the base metals in steps5,6,7,8,9 and 10 and the insoluble residue being processed to furtherconcentrate and recover the platinum group metals. The slurry exitingthe autoclave leaching stage is a finely dispersed product and is thusnot ideal for thickening and filtration. Larox type filters have beenfound to be suitable for handling slurries of this sort owing to theircompactness and possibility to conduct effective cake washing and dryingin a single stage.

Through experimentation it has also been found that in order to assistwith processing conditions downstream of the filtration stage 4 themoisture content of the cake eminating from the filtration stage shouldnot be more than 13%. Accordingly, it is advisable to increase theduration of dewatering of the material in the filter until the desirablemoisture content is achieved.

There are a number of different options which can be followed in therecovery of the base metals from the filtrate. In the embodiment of theinvention depicted in FIG. 1 the filtrate is neutralised with lime 6 toa pH of approximately 4, followed by contacting the filtrate with alime/sulphur slurry 7 at 150°C. pO₂ =1000 kPa, t=60-80 minutes, andliquid to solid ratio of 3:1 to precipitate the base metals assulphides. In effect this is autoclave leaching of the base metals.These sulphides are then recovered by flotation as a mixed Ni, Cu, Coconcentrate.

Other options which also exist for base metal recovery from the filtratewould include solvent extraction and precipitation with hydrogensulphide.

The insoluble residue 11 containing the platinum group metals emanatingfrom the filtration step 4 are passed to an oxidising roast 12 which inthe described embodiment of the invention is performed at temperaturesof 500 to 1000° C. Directly before roasting the material is mixed withlime and granulated. The addition of lime repeats the removal of sulphurto gasious phase 13 and the granulated material limits dust removal fromthe furnace. It is proposed to use a shaft furnace with the adjustmentof heating mode by heating gases obtained by burning liquid or gas fuel.

Through experimentation it has been found that the oxidising roastresults in approximately 85 to 93% of the Osmium present in theinsoluble residue being removed to the gas phase. It was also found thatalong with the Osmium about 5% of Ruthenium passes to the gas phase. Therecovery of Osmium is achieved in a scrubbing system by abadsorption.Gas eminating from the roasting stage containing sulphur and sulphuricanhydride in addition to Osmium tetraoxide is spread by recyclingsolutions in the absorbers. In this way the Osmium tetraoxide andsulphuric anhydride are removed from the gas. It is known to recoverOsmium from the off-gas of a smelter according to known processes forthe extraction of platinum group metals from a flotation concentrate.The advantage of the process of the invention over and above the knownsmelter process is that the volume of off-gas leaving the roaster issignificantly less than from a smelter which allows for improvedrecovery of Osmium in the scrubbing process.

This oxidation roast produces calcines 14 which are chlorine leached attemperatures of 20 to 90° C. in step 15. A two stage chlorination isrequired to achieve high dissolutions of Pt (in excess of 96%) and Pd(in excess of 99%) from the calcine. In tests which were conducted onthis process by the applicant it was found that Rh dissolution was low,typically approximately 13%. Nevertheless, it was found that Rhdissolution tends to increase with both increasing roasting andchlorination temperatures.

As a result of fulfilled investigations (stages 2, 3) a technologycomprising two-stage calcination chlorination leaching with thecounter-current flow of solid and liquid phases is proposed forindustrial implementation. The process conditions (for each stage) areas follows: temperature 85-90° C.; L/S ratio—3-3.5; [HC]_(initial)=170g/l; duration 2-2.5 hours and Redox-potential 950-1050 mV. provide therecovery from the cake after AOL (%): platinum 99, palladium 92, rhodium84, ruthenium and iridium 90, gold 95. The aforementioned processparameters have been found to lead to the following percentagerecoveries of the platinum group metals.

These recoveries can be increased, for example, by increasing each stageduration to 4 hours. However, this leads to the increase of iron contentin the final solution to 20 and more g/l, that is undesirable as itinterferes with later PGM adsorption from the solution.

During the additional research of the hydrochlorination—adsorptionstages there were found two items, which have to be considered while theimplementing of the technology.

It is envisaged that in place of an oxidising roast 12 a reductive roastcould be conducted on the insoluble residue 11. A hydrocarbon sourcecould be used as a reductant, which converts the platinum group metalsto the metallic state. Such a reduction would typically be done at atemperature of 650° C. Based on tests which have been conducted by theapplicant on the method of the invention it would seem that if thecalcine is reduced, as opposed to being oxidised, lower roastingtemperatures can be used.

The roasting temperature can also be lowered by subsequently forming athermal reduction of the calcine prior to chlorination. It will beappreciated that this would introduce an additional stage into theprocess.

The chlorinated slurry emanating from the leaching step 15 is cooled andfiltered 16. The filter cake is washed before disposal 17 of the residuewhich comprises the filter cake. The filtrate 18 from the filtrationstep is passed to an ion exchange adsorption unit 19 for extraction ofthe platinum group metals from the filtrate by adsorption onto ionexchange resins which are selective for platinum group metals, forexample proprietary resins such as ROSSION 11 and ROSSION 70.

From the ion exchange adsorption stage 19 the resin onto which theplatinum group metals have been adsorbed is passed through an ionitewashing unit 20 before the resin 21 is passed to a desorption unit 22.Desorption of the platinum group metals is done with thiourea accordingto known technology as is depicted diagrammatically in unit operations24, 25, 26 and 28 in the accompanying drawing. The use of thiourea mayequally be replaced with another appropriately selected desorptionchemical due to potential carcinogenic effects of thiourea.

An alternative to the fairly complex desorption stage 22 would be toburn the resin. Burning of the resin has environmental implications, butwould result in a product containing approximately 80% platinum groupmetals in an unrefined state.

In the first embodiment of the invention the platinum group metals arestripped from the resin and then either precipitated, to form aconcentrate 27 which can be further refined to the individual metal (Pt,Pd, Rh, Ru, Ir) sponges or salts.

The solution 29 from the washing unit 20 is passed through an ironextractor 30 to obtain a solution 31 that is fed to washing unit 16. Anorganic phase from iron extractor 30 is directed to an iron strippingprocess 32 to obtain a solution 33 which is fed to an iron precipitationtank 34.

FIG. 4 of the accompanying drawings depicts an alternative embodiment ofthe invention. In this embodiment the essence of the invention, namelythe three steps of base metal recovery in a leach 50, roasting 52 toconvert the platinum group metals to a form that dissolves inchlorine/HCl and the chlorine/HCl leach 54 that provides the platinumgroup metals in solution, are retained with changes to the ancillaryfeatures of the invention.

The most notable differences between the process proposed in thisembodiment of the invention and that proposed above with reference toFIGS. 1, 2 and 3 is that the conditions of the pressure oxidativeleaching of the base metals and sulphides 50 are set such that theydissolve as much of the base metals and sulphides as possible. Thisreduces the amount of Fe remaining in the solid phase, dissolvingdownstream in the HCl/Cl₂ leach of calcine and interfering with theion-exchange recovery 60 of platinum group metals. It is thereforedesirable to dissolve most of the iron during the pressure oxidativeleach step 50, followed by a separation step 56 involving pressureoxidation to precipitate iron as haematite and thereby separate it fromthe dissolved copper and nickel. Iron is then removed from the dissolvedcopper and nickel by counter-current washing or filtration, and thecopper and nickel recovered by precipitation as a bulk concentrate or bysolvent extraction.

APPENDIX PART 1 1. The PGMs concentrate Amount - 10,000 t/yr. Theconcentrate composition: Element Content, % Amount, tons ElementContent, % Amount, tons Ni 1.04 104 Ti 0.19 19 Cu 0.62 62 Pt 0.0156 1.56Co 0.023 2.3 Pd 0.00747 0.747 Fe 7.9 790 Rh 0.00263 0.263 S 1.72 172 Ru0.00527 0.527 Mg 9.3 930 Ir 0.000059 0.0059 Ca 2.3 230 Au 0.0001910.0191 Cr 2.8 280 Os 0.00012 0.012 SjO2 42.24 4224 Al 3.25 325  2. Theslurry preparation: Liquid: solid ratio = 2 The slurry is prepared inthe reactor equipped with a stirrer, V - 4m³ material - steel 3.  3. Theautoclave oxidising leaching: Process parameters: Temperature - 145 ± 5°C. Duration -2 hours Partial oxygen pressure - 0.5 MPa Total pressure -1.0-1.1 MPa The autoclave needed the number of the sections - 4, thenumber of stirrers - 4, V = 10m³. Extraction into solution, in %: Ni -75.0; Cu - 90; Co - 70; Fe - 40; Ca - 10.8; Mg - 7.25; S - 70.0 Oxygenconsumption - 300 t/yr. Sulphuric acid consumption - 900 t/y.  4.Filtration and washing Filtration rate - 0.3 m³/m²*hour Equipment - afilter with filtering area S = 10m² Insoluble residue is washed in thefiltering area. Water consumption - 0.5 m³/t of the residue. Output ofthe insoluble residue - 101% of initial.  5. Solution (filtrate) afterautoclave leaching: Amount - 22 500m³/yr. Composition, g/dm³: Ni - 3.47;Cu - 2.49; Co - 0.063; Fe_(total) - 14.04; Ca - 1.1; Mg - 3.0; H₂SO₄ -26.7  6. Neutralisation is carried out with lime (pH is adjusted up to4) at 60-70° C. for 45 minutes. Lime consumption - 1442 t/yr. Limeactivity - 90% The lime slurry is prepared in a reactor (V - 0.2 m³);water consumption - 1442 m³/yr. PART 2  7. Precipitation is carried outwith lime-sulphur slurry (LSS) at 90° C. for 45 minutes. Extractiondegree for nickel is 94-95%; copper - 99.5%; cobalt - 90%. For obtainingof the LSS, the slurry with S: CaO: H₂O ratio = 2:1:8 is heated up to95° C. and maintained for 1 hour. Liquid phase of the LSS contains(S_(mooo) S_(thio)) ˜.75 g/dm³. To prepare the LSS a 0.25 m³ reactor isrequired. Precipitation of sulphides is conducted in reactors with totalvolume of 5m³. Amount of solid phase - 3645 t/yr. The solid phasecomposition, %: Ni - 2.0 ± 0.02; Cu - 1.52 ± 0.02; Co - 0.04 ± 0.002;Fe - 8.3 ± 0.1; S - 3.45 ± 0.2; Ca - 17.2 ± 0.2. Volume of liquidphase - 24842 m³/yr. The liquid phase composition, g/dm³: Ni - 0.18;Co - 0.006; Fe_(total) - 0.41; Ca - 1.1; Mg - 2.72; H₂SO₄ - 1.8.Reagents consumption for the LSS preparation: Lime (activity - 60%) -112.5 t/y Sulphur - 225 t/y H₂O - 900 m³.  8. Flotation of slurry isaccomplished in a flotation machine according to the scheme: basicflotation and retreatment of tailings. The performance of a flotationmachine is 3.5 ± 0.25 m³ of slurry per hour.  9. Tailings afterflotation of non-ferrous metals contain: The solid phase - 3115 t/y;liquid phase - 24313 m³. The solid phase contains, %: Ni - 0.1; Cu -0.04; Co - 0.001; Fe - 6.7; Ca - 19.8; S - 0.22 The liquid phasecontains, %: Ni - 0.22; Co - 0.002; Fe - 0.41; H₂SO₄ - 1.8; Ca - 1.1; Mg-2.74. The tailings can be directed to the deposit area or afterconcentration up to liquid: solid ratio = 1:1, the concentrated slurrycan be directed to the deposit area, while solution (21198 m³/yr) - tothe oxidising leaching. 10. Non-ferrous metals concentrate (amount - 529t/y, moisture - 50%) containing, %: Ni - 13.2; Cu - 10.4; Co -0.27; Fe -18.5; Ca - 1.9; S - 22.5, is directed to the processing for extractionof non-ferrous metals into a commercial product. 11. Insoluble residueafter leaching. Moisture - 20%; amount - 10100 t/y (dry weight). Theresidue composition, %: Ni - 0.26; Cu - 0.06; Co - 0.006; Fe - 4.69;Ca - 2.03; Mg - 8.54; S - 0.51. Platinum group metals do not actuallypass into solution during leaching. 12. Oxidise roasting. Oxidisingtemperature - 1000 ± 50° C. Duration - 2 hours. A tube furnace isrequired: diam. - 1.2 m, length - 22 m. The furnace rotation speed - ω0.6 rpm. Electric motor capacity - W - 50 kW. 13. The off-gases (fromthe tube furnace) containing osmium are directed to scrubbing with thefollowing osmium recovering into a commercial product by the knownmethods. 14. The roasted material after cooling up to 60-80° C. isdirected to leaching for PGMs to be transferred into solution. Theroasted material yield is 100 ± 2% of the charge. 15. Leaching of thePGMs PART 3 For the PGMs extraction into solution the roasted materialis subject to countercurrent two-stage leaching. The process parameters:Temperature - 80° C.; Duration of each stage - 8 hours; solid: liquidratio 1:1.5 The recycled solution of hydrochloric acid (3.5N) isutilised for leaching. Chlorine consumption is 5 t/y. Leaching iscarried out in the reactors equipped with stirrers. Total volume of thereactors - 50 m³. Material - titanium alloys. Extraction from theroasted material into solution, %: Pt - 95.5; Pd - 98.0; Rh - 46.0; Au -90.0; Ru - 65.0; Ir - 70.0; Fe - 9.3; Cu - 20.0; Ni - 20.0; Co - 6.0;Al - 8.0; Ca - 20.0. 16. Filtration and washing. Filtrationtemperature - 20-30° C. Filtration rate - 1.5 m³/m^(2˜) hour Filteringarea - S = 2-2.5 m². Washing of residue is conducted with the recycledsolution of HCl (3.5N). The solution consumption - 1 m³/t of theresidue. 17. Output of washed residue (dry weight) is 98 ± 1% of theroasted material. Moisture of the residue - 20%. The residue (in amountof 9898 t/y) is directed to the deposit area. The residue composition,%: Ni - 0.21; Cu - 0.048; Fe - 4.34; Al - 3.02; Ca - 1.66; Mg - 8.71;Ti - 0.19; Co - 0.006; SiO₂ 42.68; S -0.01; Pt - 0.0007; Pd - 0.00015;Rh - 0.0014; Au - 0.000019; Ru - 0.0019; Ir - 0.000018. 18. Filtrate andsluice water are directed to PGMs sorption: The amount of the solution -23000 m³/yr. The solution composition, mg/dm³: Ni - 226; Cu - 52.5; Fe -1913; Al. - 1130; Ca - 1763; Pt - 64.78; Pd - 31.83; Rh - 5.26; Au -0.75; Ru - 14.9; Ir- 0.178. 19. Sorption of the PGMs is accomplished inthree sorption columns (two of them are used for sorption the PGMs, thethird one for desorption of the PGMs and washing). Ionite Rossion 11 isused as a sorbent. The sorbent capacity is 60 kg of the PGMs per 1 tonof the ionite. The ionite swelling factor - 3.0. Solution flow ratethrough the sorption columns 3 m³/hour. Three-column plant isacceptable; diam - 1.0 m; length - 5 (for each column). Material -titanium. 20. Washing of the ionite is carried out by water. Waterconsumption - 300 m³/yr. Washing water together with the solution aredirected to iron extraction. 21. Extraction degree during the operationis, %: Pt - 92.31; Pd - 96.85; Rh - 97.8; Au, Ru, Ir - 98.0. 22.Desorption of ionite is conducted with thiourea (C = 60 g(dm³) at 80° C.Thiourea consumption for this operation is 600 m³/y (losses bydecomposition - 25%). 23. Washing of the ionite after desorption iscarried out with water in the amount of 200 m³/yr. Eluate (in the amountof 800 m³/y) is sent to the PGMs precipitation, while washed ionite isrecycled to sorption. 24. The PGMs precipitation is carried out inreactors equipped with Stirrers. The PGMs solution is mixed with causticsolution. PART 4 Eluate composition, g/dm³: Pt - 1862; Pd- 915; Rh -15.1; Au- 2.15; Ru - 42.9; Ir - 0.516. 25. The PGMs are extracted fromeluate solution by hydrolysis at ambient temperature and pH value of 11adjusted by feeding of NaOH. NaOH consumption is 5 t/y. Eluatecontaining the PGMs is mixed with NaOH and maintained for 0.5 hour inreactor, then, while being maintained in a thickener for 20 hours, solidPGMs compounds are generating. 26. A 2.5-in diam pulp thickener isrequired to concentrate the PGMs slurry. The PGMs slurry filtration rateis 0.2 m³/m²* hour. Filtering area - 5m². 27. The PGMs concentrate inthe amount of - 3 400 kg/y, containing, %: Pt - 43.8; Pd- 21.5; Rh -3.55; Au- 0.5; Ru - 10.0; Ir- 0.12; S - 9.7; OH - 10.3, is processedwith selective extraction of PGMs into a commercial product by the knownmethods. 28. Thiourea solution in the amount of 600 m³/y is mixed withhydrochloric acid (HCl consumption - 2.6 t/y) and recycled fordesorption, while the solution in the amount of 200 m³/y is evaporatedwith the following recycling of the condensate (−200 m³/yr) for theionite washing and removal of the generated salts to deposit area fordisposal. 29. The solution (in the amount of 23 300 m³/y) containing, inmg/dm³: Ni - 223; Cu - 51.52; Fe - 1888; Al - .1115; Ca - 1740; Pt -4.92; Pd, Rh, Ru, Au < 1.0 is directed to iron extraction. 30. The ironextraction is conducted by tertiary amines in kerosene (0.8M).Extraction is accomplished in 5 steps, stripping. - in 5. Working volumeof an extractor is - 6m³. The materials - titanium, plastic. Anorganic-to-aqueous volume ratio (O:A) is 1:10 for extraction and 3:1 forstripping. 31. After extraction, 3.5 N-solution of HCl is directed toleaching (15 000m³/yr) and residue washing (10 000 m³/y). The solutiondirected to washing is mixed with hydrochloric acid (HCl consumption -217 t/y) 32. Stripping of iron is conducted by water. H₂O consumption is700 m³/y. Organic phase is recycled for iron extraction whileiron-stripped solution is directed to iron precipitation. 33. Theiron-stripped solution (in the amount of 700 m³) contains, in g/dm³:Fe - 62.6; Ni - 7.42; Cu - 1.7; Al - 36.95; Ca - 57.9. The PGMs do notpass into solution and accumulate in organic phase, from which they arethen stripped by 7 N HCl solution and directed to the PGMs sorption. 34.In order to remove Fe, Ni, Cu and other elements from the scheme, thestripped solution is processed by sodium carbonate and then dischargedas slurry into the deposit area. Process parameters: Temperature -80-90° C. Duration - 2 hours pH value - 11 Na₂CO₃ consumption - 250 t/y.

It will be appreciated that the embodiments of the invention which aredescribed above with reference to the accompanying drawings are merelyillustrative of ways of putting the invention into effect and should notbe seen as limiting on the overall scope of the invention.

What is claimed is:
 1. A hydrometallurgical treatment process forextracting platinum group metals from a flotation concentrate comprisingthe steps of: leaching the flotation concentrate to dissolve base metalsulphides in the flotation concentrate and forming a filtrate and aresidue; separating the filtrate from the residue; roasting the residueto form a calcine; and chlorinating the calcine to dissolve the platinumgroup metals into solution.
 2. A hydrometallurgical treatment processaccording to claim 1 including the steps of: adsorbing the platinumgroup metals onto an ion exchange resin; and recovering the platinumgroup metals from the ion exchange resin.
 3. A hydrometallurgicaltreatment process according to claim 1, wherein the roasting stepcomprises oxidation or reduction of the residue.
 4. A hydrometallurgicaltreatment process according to claim 3, wherein the oxidation takesplace at a temperature of up to 1000° C.
 5. A hydrometallurgicaltreatment process according to claim 1, wherein the process includes thestep of recovering osmium from an off-gas from the roasting step.
 6. Ahydrometallurgical treatment process according to claim 1, wherein thechlorination step comprises countercurrent chlorination of the calcineat approximately 80° C. and 3.5N HCl.
 7. A hydrometallurgical treatmentprocess according to claim 1, wherein the separation step comprisesfiltration followed by the additional steps of neutralisation of thefiltrate; precipitation of base metal sulphides and flotation ofprecipitated sulphides into a concentrate.
 8. A hydrometallurgicaltreatment process according to claim 2, wherein adsorption of theplatinum group metals onto the ion exchange resin is followed by:desorption of the platinum group metals from the resin with thiourea atapproximately 80° C. to form a stripped resin and an eluate, followed bywater washing of the stripped resin.
 9. A hydrometallurgical treatmentprocess according to claim 8, wherein the process includes the steps of:precipitating the platinum group metals from the eluate with causticsolution.
 10. A hydrometallurgical treatment process according to claim1, wherein said leaching step comprises leaching a slurry of saidflotation concentrate under pressure in an oxygen atmosphere in anautoclave.
 11. A hydrometallurgical treatment process according to claim1, wherein said process is in the absence of matte smelting of theflotation concentrate.
 12. A hydrometallurgical treatment processaccording to claim 1, wherein said chlorinating step compriseschlorinating said calcine in the presence of HCl and chlorine todissolve said platinum group metals into solution.
 13. Ahydrometallurgical treatment process according to claim 1, wherein saidchlorinating step is an aqueous phase chlorination and converts saidplatinum group metals to soluble platinum group metal compounds, anddissolving said platinum group metal compounds.
 14. A hydrometallurgicaltreatment process according to claim 1, wherein said roasting stepcomprises an oxidation roasting step.